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By J. C. Bradbury

The Illinois State Geological Survey's program in industrial minerals is twofold-the gathering of information on mineral raw materials and the generation of data on the use of these materials. The first function is carried out by the Industrial Minerals Section and the Clay Resources and Clay Mineral Technology Section of the Geological Group, and the second is the responsibility of the Chemical Engineering Section of the Chemical Group. Gathering in formation on raw materials includes both mapping of resources and fundamental geological research. Recent and current research projects include studies relating lithologies of limestones to their physical properties, investigation of sulfur dioxide reactivity of various types of limestone, lithostratigraphic studies of limestones of Mississippian age, investigations of heavy minerals in Cretaceous, Tertiary, and Pleistocene sands, studies of the environments of clay deposition, electron microscope studies of clays and limestones, and investigation of long-term expansion in ceramic products. A relatively recent development that relates more to dissemination of information than to research is environmental geology. This affects the Survey's industrial minerals program chiefly by stimulating the study of rock and mineral resources in metropolitan areas and reporting the results for the land- use planner as well as the engineer or mineral producer. The Survey's program of research on the use of industrial minerals was expanded and diversified in 1966. Incorporating a minerals beneficiation lab, the program is designed to develop new uses for existing mineral resources and to assist producers in implementing these ideas: Projects of the minerals engineering program have included the separation of a feldspar concentrate from certain glacial outwash and river sands, recovery of kaolinite from the waste from silica sand operations, development of the use of an Illinois clay as a binder for pellets, and investigations into the use of fly ash and fluorspar mine tailings.

Jan 1, 1970

By C. J. Weiss

"A commercially available Through-the-Earth (TTE) communications system was evaluated at an underground coal mine in eastern Kentucky. This field study was conducted as part of a larger multi-site evaluation of available TTE systems to determine their operational sensitivity. Field testing is being performed along with simulation techniques developed for geophysical surveys. The results of the field test will additionally be used to determine the applicability of these simulation techniques to TTE communications. The mine detailed in this paper was idled at the time of the study, which provided the nearest practical representation of a post-event mine shutdown. For this study, the three following communication modes were tested: surface-to-underground, underground-to-surface and surface-to-surface. Standard deployments of the TTE system using the manufacturer’s recommended procedures were able to achieve clear communications between underground and surface locations. Other arrangements of the units yielded intriguing results, some of which were predicted by computer simulations, others were unexpected. Future experimentation is planned to further evaluate the observed phenomena. INTRODUCTION The ability to establish effective communication between surface and underground personnel after a mine event is essential for both mine evacuation and mine rescue efforts. Current communication technology is reliant on physical infrastructure, such as antennas, cables, and repeaters, located throughout the mine. Such systems are easily disabled during a mine event because of their vulnerability to physical damage. A true emergency communications system would have the ability to function regardless of a mine’s physical state. Through-the-Earth (TTE) wireless communications technology has the potential to provide such a tool. Wireless TTE communications are exceptionally challenging because electromagnetic (EM), or radio, signals are subject to drastic attenuation as they travel through solid strata. In addition to attenuation, EM wave propagation is directly affected by the physical and petrographic characteristics, mineral compositions, metamorphic properties, and the degree of water infiltrations in the strata [1-5]. These strata characteristics are unique to each mine thus making the design of an effective, universally deployable, and reliable TTE communications system very difficult. At present, TTE technology is in the preliminary stages of practical deployment with only a few permissible TTE systems commercially available."

Jan 1, 2015

By Silas Li, Michael Harris-Gifford, Jason Choi, William H. Hansmire

Twin-bore tunnels with a single track in each tunnel are currently the most common subway tunnel configuration. However, a large diameter single-bore tunnel with two tracks separated by a center wall could replace twin-bore tunnels and introduce potential benefits but require necessary modifications to the subway system. In particular, a nominal 40-foot-diameter single-bore tunnel can be considered in lieu of nominal 20-foot-diameter twin-bore tunnels. This paper discusses the benefits and modifications introduced by replacing twin-bore tunnels with a single-bore tunnel, with respect to various elements of the subway related to the tunnel, stations, track, and tunnel systems.

Dec 1, 2022

By Nicholas Tabor, Norbert Fuegenschuh, Richard Taylor, Jean-Marc Wehrli

The Plymouth Tunnel forms part of the 16-mile long route of the Purple Line Light Rail Project promoted by the Maryland Transit Administration. The tunnel is approximately 1,220 feet in length with a shallow vertical alignment located beneath a residential area. Challenging mixed ground conditions included surficial soils, decomposed Saprolite, and highly weathered to fresh rock requiring the use of the Sequential Excavation Method (SEM) for construction. This paper discusses innovative approaches implemented during construction as well as challenges faced during planning and installation of a spray-applied permanent concrete liner.

Jun 13, 2021

By Chris Roos

The mining industry has made strides in planning for operations to be safe for the workforce and public, to protect the natural environment, and to maximize value for shareholders. However, many of the techniques used to accomplish these goals are based on estimated models of mineral content, production efficiency, and economic performance. The industry has begun to recognize the need for understanding the uncertainty in these models, how it affects their decisions, and what the potential impacts are for each project. The challenge has become how technical professionals can effectively communicate the impact of this uncertainty and its importance to both technical and non-technical audiences. This paper presents mining engineers (and other technical professionals) with traditional and modern principles of data visualization and technical communication that can assist with communicating the effects of uncertainty in the mine planning process.

Jun 25, 2023

By Miguel Carrera, Jose Arnal, Julian Ortiz, Maarten Haest, Fouad Faraj

ShovelSense is a complete solution for bulk ore sorting at the mine face that combines robust shovel bucket mounted X-ray fluorescence sensors, with real-time material classification, fleet management integration and supporting cloud-based data infrastructure, to ensure each truck always goes to the right destination. Correct truck destinations are a challenge in many current mining settings because the material classification of a truck, which is the key input into the truck destination, is often based on estimates at block resolution (~20x20x15m), with limited adjustment for blast movement and mixing. In this contribution, we present a geostatistical, data driven approach to estimate the potential value of bulk ore sorting at the mine-face for seven open pit mines. The approach is premised on the assumption that truck loads within blocks follow a simple gamma distribution. Standard ore control data from the seven mines have been used to estimate the potential bulk sorting value at these operations and ShovelSense-integrated dispatch datasets have been used to validate these estimates for specific benches on which ShovelSense operated. ShovelSense results validate the ore control data-based value predictions and show how at mine-face bulk ore sorting value varies between distinct deposit types. In general, deposit types that have higher bulk ore grade contrasts, display more potential ShovelSense value. The manuscript concludes with the limitations and opportunities of applying the proposed value estimation approach to evaluate at mine-face bulk ore sorting opportunities for open pit operations.

Jun 25, 2023

By Marline J. Veesaert

Urban quarries offer one of the greatest opportunities for profit in non-metallic mining with limited financial risk IF properly evaluated, planned and developed to meet market needs. Location, with respect to market, is the prime consideration. Geologic considerations and quality of rock are nearly as important. Cost factors and return on investment (ROI) are the principal financial considerations. Risk of capital is determined by weighing all factors that should be considered and how well judgements are based on information acquired. It is a simple, true axiom of business that an investment must be founded on amount of capital required with its expected return plus profit, expertise required to operate and time involved. Urban quarries generally offer excellent stone product market opportunities. Costs of land with mineral containment and processing operations are much greater than in rural areas. Environmental problems have become increasingly severe. Competition from other quarries in the market area is usually directly proportional to the profit attractiveness. Risk of capital (initial, developmental and operational) can be evaluated and reduced only by thorough analysis of the entire situation. Time and money expended on planning and preliminary investigations probably represent the most productive investment of available company resources.

Jan 1, 1977

By G. E. Erdman

This paper presents talc milling guided by a theme of "So You Want to Be In The Talc Business". Such a theme strikes a cautionary note to warn the uninitiated of potential pitfalls, and this paper will cover briefly some of the technical traps as well as one of the larger talc milling operations. It is important to our purpose to pre- sent some of the market and mineral complications that make the business of producing and selling industrial talcs much more intricate than an outsider would perceive. Metal markets by comparison are great worldwide pools into which production streams of generally identical metals can flow over a weir of market prices. Industrial talc markets are not as well defined and often must be developed by the producer working closely with a specific customer or groups of customers to prove out the attributes of the specific product from a new property. All the negatives of a new market seem to work against the introductory talc in the forms of industry or customer reluctance to change, sales prices, cost of delivery, mode of transportation, health considerations of the minerals and many more. Just locating the potential customers for a new industrial talc may be a substantial undertaking. Thus, an analogy of the worldwide metal market might compare starting a new talc operation to trying to gather up enough fallen rain drops to be able to start a production stream of a worthwhile size.

Jan 1, 1984

By J. H. Rowland, R. A. Thomas, III, C. D. Litton

Atmospheric monitoring systems (AMS) are required when using air from conveyor belt entries to ventilate working sections in U.S. underground coal mines. AMS technology has the potential to increase fire safety mine-wide, but research is needed to determine the detection and response times for fires of a variety of combustible materials. To evaluate the potential of an AMS for fire detection in other areas of a coal mine, a series of full-scale fire experiments were conducted to determine detection and response times from fires of different combustible materials that are found in U.S. underground coal mines, including high- and low-volatility coals, conveyor belts, brattice materials, different types of wood, diesel fuel, and a foam sealant. These experiments were conducted in the Safety Research Coal Mine (SRCM) of the U.S. National Institute for Occupational Safety and Health (NIOSH) located in Pittsburgh, PA, using a commercially available AMS that is typical of current technology. The results showed that through proper selection of sensors and their locations, a mine-wide AMS can provide sufficient early fire warning times and improve the health and safety of miners.

Jan 1, 2016

By W. C. Hellyer

The Granduc ore body is located beneath glacial terrain on the borders of British Columbia and Alaska, some 600 miles from Vancouver and 30 miles from Tidewater at Stewart, B. C. The concentrator and attendant facilities are located in wilderness country at an elevation of 2500 feet. The average annual snowfall is 65 feet. Surface facilities and ore body are linked by a 10 mile railroad haulage tunnel. This copper property has been developed by Granduc Operating Company on behalf of Newmont Mining Corporation and American Smelting and Refining Co. First production started in November, 1970. In August, 1966, Granduc Operating Company awarded an engineering and construction management contract to McKee-Swan Wooster, a joint venture of A. G. McKee of Canada and Swan Wooster Engineering Company Limited, to design a concentrator to treat approximately 7000 tons of ore per day. The isolated location of the Granduc ore body indicated comparatively high freight rates on consumable items of concentrator supplies such as reagents, grinding media, etc. A firm flowsheet for the flotation of the Granduc ores having been developed by the Danbury laboratory of Newmont Exploration Limited, attention was devoted to means of reducing operating costs. One obvious method was to reduce or eliminate the costs of steel grinding media by employing some form of autogenous grinding. two pilot plant operations were carried out at the facilities of Lakefield Research of Canada Limited, in which autogenous grinding of Granduc ores was investigated. The results of these investigations led to the selection

Jan 1, 1972

By R. Addison, T. Swendseid, R. Kehmeier, D. Larsen, J. Uhrie

"The development of a mining project necessarily progresses through many steps, beginning with the identification of a mineral resource and continuing through project construction. Several classifications of engineering study development are described, including preliminary economic assessment (PEA) or scoping studies, prefeasibility studies (PFS), feasibility studies (FS) and regulatory codes, each with a specific project organizational objective. This paper addresses the PEA, PFS and FS types of reports that are used by companies to evaluate projects, meet regulatory requirements and seek financing. The minimum requirements for engineering studies, focusing on the needs of financial institutions considering investment in a project, are outlined. IntroductionThe financing of mining projects in recent years has come under increased scrutiny by financial institutions owing primarily to an increasing percentage of projects whose capital expenditures had overrun project budgets or had failed to produce the economic returns anticipated during the project development process. Often, unfavorable project outcomes are not the result of widespread failures of the project engineering process but arise from inadequate engineering in limited specific areas of the project scope, such as overestimated mineral reserve, inadequate mine design, insufficient metallurgical testing or underestimated capital costs. Mill facilities and process infrastructure typically are responsible for up to 80 percent of the capital expenditures required by greenfield mining projects. They are often the project areas where capital costs are inadequately developed, and consequently are the areas highly scrutinized by financial institutions. This paper offers a perspective on the level of engineering expected by financial institutions as a prerequisite to committing support to developing mining projects.The evaluation of a mining project from exploration through development and production is a lengthy and complicated process. Mine development commitment activities for a potential project are initiated when a mineral resource is identified and continue through to the start of construction. The technical feasibility and the economic viability of each project are determined during the phases of mine development with more detailed engineering data required at each stage. There are at least four levels of engineering studies during development that are commonly acknowledged by the mining industry:"

Jan 1, 2016

By Gerald L. Colborne

The major role of financial institutions in the free world in the development of mineral resources is to provide the necessary capital in the form of loans to underwrite the development of these resources by the industry on a stable and continuous basis. This capital provided by the depositors to the bank must be invested in sound projects where the degree of risk is relatively small, the return on the investment is reasonable and the time to retire the loan is of short duration. Money is, therefore, one of the important ingredients along with technology, manpower and an exploitable resource that will provide a sound economic base in this region of development. Canadian banks are similar to banks of other western nations, although each country has some differences in their banking systems. The basic difference between the American and Canadian systems is that the former employs unit banking while the Canadian utilizes the more mobile branch system, which is prohibited or restricted in the United States and some other countries. Canadian banks are the subject of this paper, so, subsequent material will relate to Canadian banks unless indicated otherwise. It will be realized by everyone that the mineral industry is only one area of the economy in which the banks are interested and to cover the subject matter in the time allowed, the topic is restricted to mineral resource development. In Canada this is a vital sector of our economy and before we outline some of the details of financing and services involved in various aspects of mineral development, a review of the Canadian mineral industry is a necessary part of the sequence of presentation. The writer presents this material as a resource engineer advising a major bank in mineral matters, so the presentation may relate more to the financial problems of the developer

Jan 1, 1972

Horizontal stress is an important factor that contributes to roof failure during a long wall entry development, especially when the immediate roof is weak. In this study, the effects of the three factors, the stress angle between the maximum horizontal stress and entry development, the stress ratio of the maximum to the minimum horizontal stresses, and the sequence of entry development, on the stress distributions in the weak immediate roof of a three-entry system are analyzed using the finite element method, when the horizontal stress is fixed. The results have confirmed that the stress angle has a significant influence on the stress distributions. Generally, the three-entry system is in the worst stress condition when the angle is about 60°_75°. The stress ratio has little effect on the roof stresses. In addition, the sequence of entry development has also little effect on the roof stress for the specific case analyzed.

Jan 1, 1999

By M. Javier

Management is a major part of sustainability research. Many different aspects of sustainability have already received significant research attention, but the management processes in mining from the perspective of their impact on society are less understood and integrated as a holistic frame. It is evident that mining technology has received major blame in past ecological disasters and major mining-induced environmental hazards, but the management-mining process has escaped unscathed. This should not have happened. Instead, the management processes should share equal blame with obsolete technology for the failure to identify and monitor potential environmental hazards before their inception of mining. We will present an emergent style of management for the future of mining. It challenges the actual ?mind-set? of mining management. It describes balance and equilibrium: physical, chemical, and biological in nature as an understanding of nature?s language, which is Geomimic along side ?the bottom line?. The goal is to minimize loss instead of maximize profits (waste design) at the design stage of mining and to develop clean technologies (nature's values) to achieve non-social environmental impacts that provide a prosperous legacy to future generations.

Jan 1, 2010

By Mark E. Smith, Gerald J. Gierszewski

Introduction The purpose of this paper is to present a procedure used by the authors for evaluating and designing soil liner systems. This method is particularly valuable in evaluating various treatment schemes for chemically amended soil liners. A tabulation of laboratory test results on various soil types are presented to quantify the effectiveness of certain treatments. A typical liner design program includes developing and proving soil borrow sources, designing the cross-section of the liner system, developing construction specifications, and providing construction services to ensure the intended product is achieved. Material Source Development The first step in designing a soil liner is to identify and evaluate suitable borrow sources within an economical haulage range. This is best done in a two step approach: a reconnaissance level investigation to identify target areas and a detailed evaluation of those targets. Reconnaissance: The goal of the preliminary investigation is to locate potential borrow sources for liner quality soils. This includes all natural materials which can be compacted, chemically treated, or otherwise amended to yield an installed permeability at or below some target value. This requires utilization of all available data sources: Soil Conservation Service, BLM, aerial photos, USGS geologic maps, and project geologist records. The goal at this stage is to locate shallow deposits of favorable soil types. The Unified Soil Classification System provides an excellent first pass grouping. Clays, clayey sands and silts are the most favorable soil types, although silty sands and occasionally clayey gravels can make excellent liners, and are often amenable to chemical modification. The lowest permeabilities are generally achieved with CH, CL and MH soils. Once preliminary targets have been identified using visual examination, laboratory classification tests should be performed to further refine the selection. Testing at this stage should include gradation, plasticity and hydrometer analyses. Additionally, "preg-rob" testing should be done as early as practical. Preg-rob is a phenomenon where gold or silver ions in solution associate with the clay, or other, minerals. When this occurs, a portion of the gold or silver leached from the ore is actually tied-up by the clay and thus a reduces recovery. Testing for this consists of agitating a small sample of the soil in a solution containing dissolved gold or silver, preferably of similar chemical make-up as the solution which will contact the actual liner. The solution and soil are assayed before and after agitation to determine loss to the clay. A reliable estimate of the hydraulic conductivity, commonly referred to as permeability, can be developed from the D10 value by the use of Hezen's formula: K = 100 (D1012 This relationship is limited to soils where the finer particles do not move due to the force of flowing water (i.e.: "hydrodynamic stabilitym)(1). Additionally, the effect of platty particles on permeability is not as predictable as the effect of equidimensional particles. D10 is the sieve opening size at which 10% of the material is finer. Plasticity is also important from several standpoints. Constructability is directly related to plasticity. Very plastic clays and non-plastic silts both tend to be difficult soils, while medium plastic clays and clayey sands are generally very desirable. Post construction performance is also related to plasticity (e.g. swelling, shrinkage cracking, frost heave, etc.). Additionally, low plasticity silts and silty sands generally do not respond well to chemical amendment. Source Development: The result of the reconnaissance evaluation should be an estimate of the relative probability of developing a suitable borrow source within an economical haul distance. Of course, "economical distance" depends on the degree of handling and treatment the borrow material requires, as well as the cost of synthetic alternatives. The purpose of the detailed investigation is to prove out quantity and quality of material sources, and determine design parameters such as degree of compaction, mixing, treatment and thickness of liner. The emphasis of the testing program should be permeability and strength. Strength becomes increasingly important as the slope of the liner and the height of the heap increase. Permeability testing should evaluate the effects of compaction, water content, mixing and chemical treatment where appropriate. The effects compaction and water content during compaction have on

Jan 1, 1987

By LeEtta M. Shaffner, John R. Marks

INTRODUCTION In the last twelve years, underground stoping methods at the Homestake Gold Mine have evolved from open cut-and-fill with jacklegs, cribbed raises and electric slushers to ramp-based mechanized cut-and-fill (MCF) and vertical crater retreat (VCR) with dieselpowered loaders, trucks and drill jumbos. The evolution was swift. Unfortunately, ventilation practices have had trouble keeping pace. Stope ventilation is still too dependent on auxiliary fans and coolers. A single large MCF stope might contain up to eight headings, each of which requires significant ventilation resources when a diesel loader is present. Air is shortcircuited all too frequently through mined-out VCR panels. These factors and the recent rapid decent of the center of mining prompted a study to improve ventilation practices and to more accurately project future requirements. BACKGROUND In 1992, the underground portion of the Homestake Gold Mine produced 8336 kg of gold from 1407 ktons of ore (268,000 oz from 1,551,000 st). The ore is hosted in low to medium grade meta-sediments. Stoping took place from 420 to 2347 m below the surface (1400 to 7700-ft levels). The weighted center of mining was 1662 m deep (5450-ft level) in 1990 and is projected to be 1890 m (6200-ft) in 1993. The deepest level is 2440 m (8000-ft) where the virgin rock temperature (VRT) is 56.1°C (133F). The mine is ventilated by 504 m3/s (1,069,000 cfm) measured at mid-exhaust-circuit density. The air-conditioning system includes an 8.1 MWR (2300 ton) controlled recirculation plant, a 2.0 MWR (580 ton) chilled water plant, a 1.0 MWR (290 ton) exploration drift refrigeration plant, 28 spot-coolers totaling 3.4 MWR (960 tons) and 35 spray coolers totaling 1.5 MWR (420 tons). The mine employs 117 diesel units with a total nameplate rating of 7961 kW (10,672 hp). These units include twenty-four 1.5m3 (2-yd) loaders, twenty-six 2.7m3 (3.5-yd) loaders, fourteen 3.8m3 (5-yd) and two 7.6m3 (10-yd) trucks, and assorted utility vehicles and drill jumbos. THE STUDY In April 1991, the University of Nevada-Reno (Mackay School of Mines) and the South Dakota School of Mines and Technology cooperated with Homestake on an MCF study. Mackay instrumented a stope with thermocouples, air velocity meters and hygrometers (Duckworth, 1992). South Dakota Tech conducted a finite-elements computer analysis of a back-filled MCF stope with/without light-weight shotcrete insulation on the sidewalls and back (Chellam, 1992). Homestake conducted an empirical analysis of deep-level MCF stoping. This paper describes the Homestake study. Twenty-three of the forty MCF stopes deeper than 1800 m were surveyed at least once during the last quarter of 1992. The stopes not included were being cable-bolted or back-filled. Figure 1 shows the complexity of one of the ramp-based stoping areas included in the survey. This particular stope has six separate production headings. A baseline wallrock heat load was derived for each stope or heading from psychrometric calculations. Broken ore, waste rock fill and fissure water were noted when present and the effects included in the baseline heat load. Other heat sources often mentioned in the literature such as metabolic heat from workers, heat from explosives and small electric loads were neglected. Fan heat was considered part of the ramp & crosscut heat load and thus not included in the study. Diesel heat was addressed separately. RESULTS Survey results, plotted as the heat flux against VRT, are shown in Figure 2. The equation for the regression line is: W/m2 = 2.1236*VRT - 75.405 The 0.31 correlation coefficient is poor which implies that the results should be used cautiously. Previous experience strongly suggests that differences in productivity are most likely responsible for the scatter in data points. A rapidly advancing stope will have a

Jan 1, 1993

By P. R. Taylor

The ability to process complex precious metal sulfide concentrates containing antimony and arsenic in an economic and environmentally safe manner is an important problem facing the mining industry in the United States. Preliminary experiments have been performed to evaluate the possibility of treating a complex sulfide concentrate to recover antimony and precious metals while fixing the arsenic and sulfur as calcium arsenate and calcium sulfate. The process involves: • the oxidation of the antimony and arsenic minerals in the presence of soda ash (to collect the sulfur dioxide and to keep the antimony and arsenic from forming volatile compounds), •the selective leaching of the arsenic from the roast calcine with a heated caustic solution, •the precipitation of the arsenic and sulfate with calcium hydroxide and/or calcium chloride, •the reduction of the residual antimony (collecting the precious metals), •fuming of the metal to form a marketable antimony oxide, and •crucible reduction of the residue to obtain a gold and silver dore. Batch experiments, using synthetic concentrates of stibnite and arsenopyrite, were performed to evaluate the first stage of this process. Variables studied included mixture weight fractions, temperature, gas flow rate, oxygen partial pressure, and time. All of the off gas dust and sulfur containing gases were collected and analyzed. The roast solid residues were subjected to arsenic leach tests, and the leach solutions were subjected to arsenic and sulfate precipitation tests. Introduction The ability to treat a precious metal ore containing appreciable amounts of antimony and arsenic is dependent upon the characteristics of the ore. For some ores, selective flotation can be used to separate the antimony and arsenic minerals, which are then treated separately. Unfortunately, many ores are not treatable in this way, and the concentrates obtained may contain appreciable contamination (which limits the ability to make acceptable grade antimony products). If we consider the added value of the precious metals in the concentrate, then it is worthwhile to evaluate the potential for new methods to treat these types of concentrates. In addition, the environmental difficulties associated with arsenic disposal must be taken into account in any proposed process. If there is appreciable arsenic present, then it must be either recovered as a marketable product or placed in an insoluble form. The problem addressed in this research program has become increasingly important as the precious metals mining industry is encountering more ores of this type. A large deposit of this type exists in Idaho where the antimony occurs primarily as stibnite, with minor amounts of kermesite and tetrahedrite (Cooper, 1951). The gold appears in association with arsenopyrite with minor amounts of other arsenic minerals. Flotation has been used to obtain an antimony concentrate. The concentrate, however, is contaminated with appreciable amounts of arsenic and also contains precious metals. If the ore or concentrate contains appreciable amounts of base metals, then the recovery process becomes much more difficult. There are a number of other known deposits of this type in Alaska, Canada, Montana and elsewhere. The objective of this research program was to develop a fundamental understanding of the oxidation and leaching chemistry of the arsenic and antimony minerals so that technology may be developed to treat these ores in an economic and environmentally safe manner. Wendt (1949) and Koster and Royer (1940) discussed the possibilities of oxidative roasting of stibnite concentrates containing arsenic. Wendt describes a process used at a smelter in Czechoslovakia (in the 1930s) where roasting was carried out in rotary furnaces at 723° K (500° C). Most of the arsenic was removed as fume, as well as 25% of the antimony. No sulfur dioxide containment was described. The oxide fume from the furnace went to a heated caustic water leach, where the arsenic was preferentially dissolved and then precipitated with calcium chloride as calcium arsenate following solid-liquid separation. The antimony oxide was sold, and the solid oxide discharge from the rotary furnace was sent to a blast furnace where the antimony metal, containing the precious metals, was recovered. The metal was fumed to obtain antimony oxide, and the precious metals were left in the residue and collected with copper. The copper

Jan 1, 1994

By N. D. Jagger, I. M. Arbuthnot

Introduction Over the last five years, a large number of small- to medium-sized carbon-in-pulp treatment plants have been built in Australia, most designed to treat between 250,000 t/a and 1.5 Mt/a of ore. Because of the limited capital resources and tight cash-flow positions of these relatively small mining companies, the primary requirement was often to get a plant built and operating in a short period of time and at minimal capital cost. Therefore, since the inclusion of both pre-leach and tailings thickeners represents an obvious and significant capital cost, most of these plants were built without thickeners or even detailed, cost-benefit analyses on their inclusion. In some cases, the increasing use of High Rate Thickeners (HRTs) in the mineral processing industries has, however, resulted in a reassessment, because of their considerably lower cost. This reassessment was triggered primarily by the need to conserve water in arid mining areas where borefields are costly to install and water is limited. With the startup and operation of these installations, the resulting significant savings in cyanide consumption has been recognized, in many situations, as a primary justification for the installation of HRTs. Solution balances Degradation of cyanide occurs in the tailings water discharge to slime dams. The degree of degradation (cyanide loss) in the water recovered depends on a number of factors, but it is usually assumed to be about 90%. The most important mechanisms of CN loss are through HCN losses and oxidation by oxygen in the air, which also assists in the hydrolysis of CN. These mechanisms are supported by the large dam surface area and the long retention time of the tailings water in the dam. By thickening the CIP tailings at the plant and recovering as much tailings water as immediately possible, these losses are avoided. The retention time in an HRT is less than three hours, and the surface area is relatively small. Therefore, CN losses are negligible, which is not necessarily the case in conventionally-sized thickeners. Fig. I shows a block diagram of a 100-t/h gold plant without a thickener. In this example, 50% of the tailings water pumped to the tailings dam is recovered, and the CN concentration of the returned water is 10% of the tailings CN concentration of 150 ppm. Fig. 2 represents a plant with an HRT on tailings, thickening to 55% w/w solids. In this case, the thickened tailings are pumped to the dam, and 25% of the contained water is recovered. The recovered water and the mill make-up water are not sent directly to the mill; instead, they are added to the thickener feed and mixed with it prior to thickening. By doing this, the tailings are effectively washed, and the additional cyanide is recovered. Solution balances over these two circuits show cyanide recoveries of 5% and 65%, respectively. Thus, the thickener use increases cyanide recovery by 60%. Fig. 3 shows a two-stage HRT circuit in a countercurrent decantation (CCD) configuration. In this example, an additional 13% of cyanide is recovered through the use of the second-stage unit. This configuration can be justified when residual cyanide levels are high. Capital and operating costs - Case study 1 Illustrative cost figures are based on a CIP tailings thickener installed, in early 1988, as part of Dominion Mining's treatment plant at Paddy's Flat near Meekatharra. (All costs within this paper, unless stated otherwise, are in Australian dollars.) [Assumptions: Feed rate 150 t/h Ore moisture content 5% Leach density 40% solids Underflow density 55% solids Residual cyanide in tailings 150 g/m3 Flocculant dosage 15 g/t Consumable costs: Water $ 0.40/m3 Cyanide$ 2.00/kg Flocculant$ 4.50/kg Power$ 0.12/kWh Length of tailings pipeline2,500 m Capital costs of the major items involved in the thickener installation are given below: Thickener unit, 15-m diam $ 230,000 Water return pumps 12,000 Water recirculation pumps 28,000 Feed box 5,000 Flocculant make-up system 18,000 Flocculant storage tank and dosing pumps 11,000 Piping and valves 85,000 Electrics and instruments 47,000 Civil works 29,000 Installation 69,000 Total cost$ 534,000] Savings in capital costs that can be attributed to the

Jan 1, 1993

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

The reliable quality assessment of grouting operations is one of the key issues in establishing impermeable grout curtains around shafts, drifts and tunnels. Reliable opera¬tional information on the decrease in the permeability of rock during the grouting process facilitates the assessment of the effectiveness of the completed curtain and assures production quality. Quality control and assurance also fa¬cilitates the early detection and elimination of possible de¬fects in the grout curtain. Thorough control of the grouting process provides the capability to evaluate decreases in the permeability of water-producing fracture zones prior to the beginning or the continuation of excavation of the shaft, tunnel or drift. In order to achieve this objective, post¬grouting hydrodynamic investigations are conducted prior to the initiation or re-initiation of work on the underground workings to measure the decrease in the permeability coef¬ficient of aquifers during grouting operations. 8.1 QUALITY ASSURANCE AND CONTROL OF THE ISOLATION OF ROCK BY GROUTING THROUGH HOLES DRILLED FROM THE SURFACE Quality assurance and control are based on measuring the decrease in the permeability coefficient of water-pro¬ducing horizons as a result of injecting grout into the drill¬holes. For this purpose, the value of the acceptable residual permeability coefficient is calculated beforehand based on the acceptable value of ground water inflow into the shaft, drift or tunnel as specified by the "construction norm and integrated practice (SNIP)" (Anon., 1976). Having ob¬tained the values of aquifer thickness, permeability, and head distribution for each aquifer from flowmetric data (dis¬cussed in Chapter 3), the average value of the permeability coefficient of all aquifers intersected by the shaft is calcu¬lated as [ ] where Ki is the permeability coefficient of the "ith" aqui¬fer; Pi is the pressure of the "ith" aquifer; and M; is the thickness of the "ith" aquifer. The total expected inflow of ground water into the shaft is [ ] where RK is the radius of the contour of influence; RCFB is the radius of the shaft; and p is the coefficient of dynamic viscosity of the ground water. The necessary reduction of the coefficient of permeabil¬ity of the aquifer [ ] where Q is the expected inflow of ground water into the shaft from eq. 8.2; Q.-fin is the acceptable inflow of ground water; and a is the safety factor. By knowing the average permeability value of the rock and the permeability reduction coefficient, it is possible to determine the value of the acceptable residual permeability coefficient for all the aquifers [ ] The values of the permeability coefficient, pressure, and thickness for each aquifer are determined from the data gathered during the hydrodynamic investigations conducted in grout holes prior to beginning the grouting operations as discussed in Chapter 3. After injecting the grout into the first of the holes, the injection into each of the following grout holes is preceded by repeated hydrodynamic investi¬gations. Using the data obtained for each water-producing zone, a graph is constructed where the numbers representing the grout holes in their order of grouting are placed on the abscissa, and the values of the aquifer permeability coeffi¬cient are placed on the ordinate (Fig. 103). This curve char¬acterizes the change in permeability of the given hydro¬stratigraphic unit as grouting operations are completed, one hole at a time. The results of grouting operations are considered to be positive when each sequential observation records a steady decrease in the permeability of the aquifer. The permeabil¬ity measured prior to the injection of the grout into the last hole must be lower than or equal to the acceptable perme¬ability as determined by the application of eqs. 8.1 through 8.4. Table 8.1 presents illustrative data on the hydrogeo¬logic properties of aquifers at a mine shaft prior to the initiation of grouting operations. Data also are presented on the reduction in the permeability coefficient during the in¬jection of the grout. The shaft in question intersects eleven aquifers. Based on data concerning aquifer pressure, thick¬ness, permeability, and the expected inflow rate of water for each of the strata, the average permeability coefficient is calculated according to eq. 8.1 as K = 0.2 x 10-12 m2. The

Jan 1, 1993

By Rudolf E. Greuer

Introduction The Institution of Mining and Metallurgy and the Institution of Mining Engineers organized the Third International Mine Ventilation Congress held in Harrogate, England. Sixty-one papers were presented and about 300 participants registered. About 30% of the participants came from Great Britain, 20% from South Africa, and about 5% each came from Australia, Canada, China, France, US, and West Germany. The remainder of the participants came from 18 other countries. Eleven of the papers presented dealt with methane, three with diesel exhausts, four with dust, four with radioactivity, 18 with heat, and 15 dealt with main and auxiliary ventilation. Six of the papers dealt with mine fires, which is a boundary region between main ventilation and gas concerns. Ventilation Congress proceedings are available. See ME, December 1984, page 1687, New Books page. With CH4 and radioactivity topics, the Australians dominated since they are becoming large coal and uranium producers. With diesel exhausts, the most important problem in highly mechanized mines, the North Americans were prominent. The South Africans, working in 90% SiO2 in their gold deposits, were preeminent in presentations on dust. They also led in topics related to heat, joined by the West Germans whose mines are getting deeper. In main and auxiliary ventilation, Great Britain and West Germany provided the majority of contributions. Methane The large number of methods for the precalculation of CH4 liberation in longwall mining all contain three elements: gas content of the coal, gas emitting zone (or influence zone of face), and degree of gas emission. Determination of the gas content, commonly accomplished by taking coal samples, does not pose much difficulty. But this is not the case with the other two factors. Most existing approaches rely on rock mechanics observations. Some only rely on intuition or speculation. The West German coal mines conducted a large research project between 1977 and 1982, in which gas pressures around longwall faces were measured. Gas pressures and gas contents can be related. Therefore, influence zone and degree of gas emission can be determined simultaneously. The precalculation of CH4 liberation in room-and-pillar mining is simpler. Since foot and hanging walls remain essentially intact, gas pressure distribution and gas flow can be calculated using hydraulics equations. A key to these calculations is permeability. A group of Australian researchers reported that the permeabilities of rock under three dimensional stress differs from rock under destressed conditions. This fact was known, but no systematic observations existed. These were provided. Methane drainage has been practiced for more than 40 years. It is used with great success in longwall operations. Some of the West Germans think that methane drainage from the footwall is neglected. Under certain geological conditions, they claim this is as important as drainage from the hanging wall. A research project was presented in support of this claim. Considerable efforts have existed for more than 10 years to use methane drainage in room-and-pillar mining. A paper described the accomplishments of Consolidation Coal Co. Since coal and rock are not fractured as much as in longwall mining, gas transport takes place by the slow diffusion from micropores into the cleats. Then, it is transported by laminar flow along the cleats until a drainage borehole or the mine workings are reached. Boreholes about 300 m (1000 ft) long can reduce the gas content of a band of coal 100-m (330-ft) wide by 50% in one year. A paper from India described methane drainage from gob areas. A French paper reported that especially high methane concentrations can be drained from gob areas if the face ventilation is descensional. Methane buoyancy and ventilating pressures compensate each other, and little air dilution by leakage takes place. Another French paper described a newly developed gas and air velocity monitoring system making use of a microcomputer. Four papers on methane dealt with gas outbursts. Two Australian contributions described the problems encountered with CH,-CO2 mixtures and the attempts to solve them with drainage boreholes having 10 to 20-kPa (1.5 to 3-psi) suction pressures. A Hungarian paper described problems and attempts to reduce the risks of gas outbursts in Hungary. Prediction of gas outbursts in coal seams was the focus of an English paper. Observations show that they are frequent when coal, through its geological past, has become soft and brittle with a resulting increased desorption rate. An instrument for gravimetric de-

Jan 1, 1985